CN111259542A - Method for calculating impact resistance of anchoring support of roadway roof - Google Patents
Method for calculating impact resistance of anchoring support of roadway roof Download PDFInfo
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Abstract
The invention discloses a method for calculating the impact resistance of a roadway roof anchoring support, which comprises the steps of firstly respectively calculating the energy of a roadway roof anchor rod and an anchor cable before the roadway roof anchor rod and the anchor cable generate ultimate deformation damage, then calculating the energy limit value absorbed by a roof support system before the roof support system is damaged according to the spacing, wherein the energy applied to the support system by a roof is the sum of kinetic energy and potential energy when rock burst occurs, and the energy is equal to the energy limit value of the support system when the rock burst occurs, so that the minimum speed causing the failure of the roof support system is obtained; the method has the advantages that the maximum rock mass vibration speed corresponding to the maximum mine vibration energy level obtained by mine monitoring is obtained and compared with the obtained minimum speed, the impact resistance of the current roof anchor net cable support is judged according to the comparison result, and theoretical guidance is provided for whether the subsequent support optimization needs to be carried out, so that the method has important practical significance and social benefit in the aspects of roof subsequent support, coal mine safety production, personnel safety and danger and the like, and has wide application prospect.
Description
Technical Field
The invention relates to a method for calculating the impact resistance of a roadway roof anchoring support, and belongs to the technical field of roadway support monitoring.
Background
Rock burst is a typical mine dynamic phenomenon and has great harmfulness. The dynamic phenomenon instantly releases a large amount of elastic deformation energy accumulated in the coal rock mass in a sharp and violent mode, so that the coal rock mass is damaged and generates strong vibration, the broken coal rock is thrown to a roadway excavation space by the dynamic phenomenon, and strong sound is emitted, so that equipment damage, roadway damage, casualties and the like are caused.
Along with deepening of mine mining depth, increasing of mining intensity and complexity of mining layout, stress conditions of surrounding rock mining are gradually worsened, and rock burst disasters caused by the worsening of the stress conditions are increased and strengthened rapidly. In the prior mine with the over-impact ground pressure, the disaster of the over-impact ground pressure is more serious, and the over-impact ground pressure gradually begins to appear in the mine which has not occurred before. Therefore, the capability of resisting rock pressure needs to be considered when the anchor net cable is used for supporting the top plate. Therefore, how to accurately calculate the impact resistance after the roof anchor net cable is supported is very important for the safety of mine personnel and the safety production of mines.
The impact ground pressure resistance of the supported top plate is not accurately calculated after the top plate is supported by the existing anchor net cables, so that the support cannot be timely strengthened at the part with weak impact resistance, and finally the roadway is damaged after the impact ground pressure occurs; therefore, how to accurately calculate the impact resistance after the roof anchor net cable is supported is the research direction of the industry.
Disclosure of Invention
Aiming at the problems in the prior art, the invention provides a method for calculating the impact resistance of a roadway roof anchoring support, which can accurately obtain the impact resistance of a roof anchor net cable after support, thereby effectively ensuring the safety of mine personnel and the safety production of a mine.
In order to achieve the purpose, the invention adopts the technical scheme that: a method for calculating the impact resistance of a roadway roof anchoring support comprises the following specific steps:
i, confirm the anchoring section length of tunnel roof stock: firstly, acquiring the compression coefficient of the used anchoring agent, the length of the anchoring agent in an anchor drilling hole, the diameter of the anchoring agent in the anchor drilling hole, the diameter of the anchor drilling hole and the diameter of an anchor rod; the obtained data are then substituted into the following formula:
in the formula, L1The length of the anchoring section of the anchor rod is expressed in mm; k is a radical ofSolid 1Represents the compressibility of the anchoring agent in units of 1; l isSolid 1The length of the anchoring agent in the anchor rod drilling hole is shown, and the unit is mm; dSolid 1The diameter of the anchoring agent in the anchor rod drill hole is shown, and the unit is mm; dDrill 1The diameter of the anchor rod drill hole is shown, and the unit is mm; dRodThe diameter of the anchor rod is expressed in mm;
II, determining the ultimate elongation of the roadway roof anchor rod, and then calculating the energy consumed before the anchor rod generates ultimate deformation and is broken according to the ultimate elongation of the roadway roof anchor rod: firstly, acquiring the non-anchoring section length of the anchor rod and the elongation of the anchor rod, and substituting the acquired data into the following formula:
he1＝Lis not 1×μ1
In the formula, he1The ultimate elongation of the anchor rod is expressed in m; l isIs not 1The length of the non-anchoring section of the anchor rod is expressed in m; mu.s1Indicating the elongation of the anchor rod;
calculating the energy consumed before the anchor rod generates the limit deformation and is broken according to the obtained limit elongation of the anchor rod;
III, determining the length of an anchoring section of the roadway roof anchor cable: firstly, acquiring the compression coefficient of the used anchoring agent, the length of the anchoring agent in an anchor cable drilling hole, the diameter of the anchoring agent in the anchor cable drilling hole, the diameter of an anchor cable drilling hole and the diameter of an anchor cable; the obtained data are then substituted into the following formula:
in the formula: l is2The length of the anchorage section of the anchor cable is expressed in mm; k is a radical ofSolid 2Represents the compressibility of the anchoring agent in units of 1; l isSolid 2The length of the anchoring agent in the anchor cable drill hole is shown, and the unit is mm; dSolid 2The diameter of the anchoring agent in the anchor cable drill hole is shown, and the unit is mm; dDrill 2The diameter of the anchor cable drill hole is shown, and the unit is mm; dCableThe diameter of the anchor cable is expressed in mm;
IV, determining the ultimate elongation of the roadway roof anchor rod, and then calculating the energy consumed before the anchor cable is subjected to ultimate deformation and breakage according to the ultimate elongation of the roadway roof anchor cable: firstly, acquiring the non-anchoring section length of the anchor cable and the elongation of the anchor cable, and substituting the acquired data into the following formula:
he2＝Lis different from 2×μ2
In the formula, he2The ultimate elongation of the anchor rod is expressed in m; l isIs different from 2The length of a non-anchoring section of the anchor cable is expressed in m; mu.s2Representing the elongation of the anchor cable;
calculating the energy consumed before the anchor cable is broken due to the extreme deformation according to the obtained extreme elongation of the anchor cable;
and V, determining the energy limit value absorbed by the roof support system before the roof support system is broken according to the arrangement distance condition of roof anchor rods and anchor cables, wherein the energy limit value is as follows:
Edz＝ρ1×Eg×η1+ρ2×Eds×η2
in the formula, EdzRepresenting the limit of energy absorbed by the roof support system before rupture, in kJ/m2；ρ1The density of the anchor rod arranged on the top plate is shown, and the unit is root/m2；η1The unbalance coefficient of the energy consumption of the broken anchor rod is represented, and the value is 0.8; rho2Indicating anchor cableDensity in the roof arrangement, in roots/m2；η2The imbalance coefficient of energy consumption when the anchor cable is broken is represented, and the value is 0.9;
VI, in a roof supporting system with rock burst, kinetic energy transferred to the roof rock mass and potential energy of sinking of the roof rock mass are transferred to the supporting system to become elastic energy of the supporting system, and in a critical state, the elastic energy borne by the supporting system is equal to the energy limit value of the supporting system, and the minimum speed v causing the roof supporting system to lose efficacy is calculated: the elastic energy borne by the supporting system is equal to the kinetic energy transferred into the roof rock mass and the sinking potential energy of the roof rock mass, and the specific formula is as follows:
in the formula, EdThe kinetic energy of the surrounding rock of the roadway in unit area after the rock burst occurs is expressed in kJ/m2；EhThe unit area potential energy applied to the supporting system by the roof coal blocks due to impact glide is expressed in kJ/m2(ii) a m represents the mass of the rock mass of the surrounding rock of the roadway participating in the impact process, and the unit is kg; v represents the impact velocity of the surrounding rock of the roadway, and the unit is m/s; s represents the cross-sectional area of the roadway participating in the impact process, and the unit is m2(ii) a g represents the gravity acceleration of the position where the roadway is located and the unit is m/s2(ii) a Δ h represents the distance of the top plate coal block sliding down due to impact, and the unit is m; wherein S, m, g and delta h are known data;
due to the critical state, the elastic energy to which the timbering system is subjected is equal to the energy limit of the timbering system, i.e. EdzE; the minimum velocity v that causes the roof support system to fail is given by:
thereby deriving a minimum velocity v that causes failure of the roof support system;
VII, evaluating the impact resistance of the roadway roof anchoring support: acquiring data of the mine earthquake energy level obtained by monitoring the mine, the historical mine earthquake energy level or the surrounding mine earthquake energy level, selecting the maximum rock mass vibration speed corresponding to the maximum mine earthquake energy level to compare with the minimum speed v obtained in the step VI, and if the maximum rock mass vibration speed is less than the minimum speed v, indicating that the impact resistance of the roadway roof anchoring support meets the required requirement; and if the maximum vibration speed of the rock mass is greater than or equal to the minimum speed v, the requirement on the impact resistance of the roadway roof anchoring support is not met, and at the moment, subsequent support reinforcement work is carried out according to the difference value between the maximum vibration speed of the rock mass and the minimum speed v.
Further, the energy E consumed before the anchor rod generates the limit deformation and is broken is calculated in the step IIgThe following formula:
Eg＝Ef1×he1×f1
in the formula: ef1The breaking load when the anchor rod generates ultimate deformation and is broken is expressed, and the unit is kN/root; h ise1The ultimate elongation of the anchor rod is expressed in m; f1 represents the elongation safety factor when the anchor rod is in tensile failure; wherein Ef1And f1 are known values.
Further, calculating energy E consumed before the anchor cable is subjected to ultimate deformation and breakage in the step IVdsThe following formula:
Eds＝Ef2×he2×f2
in the formula: ef2Representing the breaking load when the anchor cable is subjected to ultimate deformation and breaking, wherein the unit is kN/root; h ise2The ultimate elongation of the anchor cable is expressed in m; f2 represents the elongation safety factor when the anchor cable is damaged by stretching; wherein EdsAnd f2 are known values.
Compared with the prior art, the method provided by the invention has the advantages that the supporting conditions of the anchor rod and the anchor cable are determined, and the energy E of the anchor rod and the anchor cable on the top plate of the roadway before the extreme deformation damage is generated is firstly and respectively calculatedgAnd EdsThen, according to the row spacing of the arrangement of the roof anchor rods and the anchor cables, calculating the energy limit value E absorbed by the roof support system before the roof support system is damageddzThe energy applied to the supporting system by the top plate is kinetic energy when rock burst occursAnd the sum of the potential energy is changed into elastic energy E of the supporting system, and when the elastic energy is in a critical state, the elastic energy is equal to the energy limit value of the supporting system, so that the minimum speed v causing the failure of the roof supporting system is comprehensively obtained; the method comprises the steps of obtaining data of the mine earthquake energy level obtained by mine monitoring, the historical mine earthquake energy level or the surrounding mine earthquake energy level, selecting the maximum rock mass vibration speed corresponding to the maximum mine earthquake energy level to compare with the obtained minimum speed v, judging the impact resistance of the current roof anchor net cable support according to the comparison result, providing theoretical guidance for subsequent support optimization, and finally effectively ensuring mine personnel safety and mine safety production.
Drawings
FIG. 1 is a computational flow diagram of the present invention.
Detailed Description
The present invention will be further explained below.
Example 1:
rock burst has been observed in a mine. The mine return airway has a potential impact hazard. After the roadway roof is supported by the anchor net cables, the impact resistance of the roadway roof is determined by the method, and the method comprises the following specific steps:
i, adopting an MSGLW500 type high-strength left-handed non-longitudinal rib threaded steel anchor rod as a top anchor rod of the return airway, wherein the diameter of the anchor rod is 20mm, the length of the anchor rod is 2400mm, the top anchor rod is lengthened and anchored by adopting 1 CK2335 and 1Z 2360 resin anchoring agent, the anchoring length is 600mm, the anchoring diameter is 23mm, and the compression coefficient of the anchoring agent is 0.8; the anchoring section length of the anchor rod which is supporting the roadway roof can be calculated:
II, knowing that the length of the anchor rod is 2400mm, so that the length of the non-anchoring section is approximate to 1947mm, the elongation of the anchor rod is 15%, and the ultimate elongation of the anchor rod is 0.292 m; for safety, taking 80% of the total elongation as the elongation at constant resistance tensile failure, the energy consumed by the roadway roof bolt before the ultimate deformation failure can be calculated:
Eg＝Ef1×he1x f1 ═ 210.5 × 0.292 × 80%
III, air return tunnel roof anchor cable adoption
The steel strand, the diameter of the anchor cable is 17.8mm, the length of the anchor cable is 6300mm, each anchor cable is lengthened and anchored by 1 CK2335 and 2-3Z 2360 resin anchoring agents, the anchoring length is 600mm, the anchoring diameter is 23mm, and the compression coefficient of the anchoring agents is 0.8; the anchoring section length of the anchor cable which is supported on the roadway roof can be calculated, and the calculation formula is as follows:
and IV, knowing that the length of the anchor cable is 6300mm, therefore, the length of the non-anchoring section is approximately 5900mm, the elongation is 3.5 percent, and the ultimate elongation is 0.207m, for safety, 80 percent of the total elongation is taken as the elongation when the constant resistance tensile failure occurs, and the energy consumed by the roadway roof anchor cable before the ultimate deformation failure occurs can be calculated:
Eds＝Ef2×he2x f 2-353 × 0.207 × 80% -58.46 kJ/root
V, the density of the roof bolt of the known return airway is 2.63 pieces/m2(ii) a The density of the anchor cables is 0.13 cable/m2The limit value of the energy absorbed by the roof support system before the rupture can be calculated:
Edz＝ρ1×Eg×η1+ρ2×Eds×η2＝2.63×49.17×0.7+0.13×58.46×0.7＝95.84kJ/m2
VI, the coal seam density of the known return airway roof is 1.4 multiplied by 103kg/m3Taking 1m in the axial direction; the kinetic energy and potential energy applied to the support system by the top plate when rock burst occurs are as follows:
due to the occurrence of impactIn the critical state of the pressed roof support system, the kinetic energy transferred into the roof rock mass and the sinking potential energy of the roof rock mass are transferred to the support system to be changed into the elastic energy of the support system, and in the critical state, the elastic energy borne by the support system is equal to the energy limit value of the support system, namely EdzE, the minimum velocity v that causes failure of the roof bracing system can therefore be found, and is calculated as:
VII, evaluating the impact resistance of the roadway roof anchoring support: the mine is used for monitoring the existing mine earthquake energy level or the historical mine earthquake energy level or the mine earthquake energy level around the mine, the mine earthquake maximum vibration speed corresponding to the mine earthquake maximum energy level is selected to be 8.45m/s, the mine earthquake maximum vibration speed is compared with the support failure minimum speed v, and the mine earthquake maximum vibration speed is smaller than the support failure minimum peak speed v, so that the roadway roof anchoring support impact resistance can meet the required requirements, and the support effect is good.
Claims (3)
1. A method for calculating the impact resistance of a roadway roof anchoring support is characterized by comprising the following specific steps:
i, confirm the anchoring section length of tunnel roof stock: firstly, acquiring the compression coefficient of the used anchoring agent, the length of the anchoring agent in an anchor drilling hole, the diameter of the anchoring agent in the anchor drilling hole, the diameter of the anchor drilling hole and the diameter of an anchor rod; the obtained data are then substituted into the following formula:
in the formula, L1The length of the anchoring section of the anchor rod is expressed in mm; k is a radical ofSolid 1Represents the compressibility of the anchoring agent in units of 1; l isSolid 1The length of the anchoring agent in the anchor rod drilling hole is shown, and the unit is mm; dSolid 1The diameter of the anchoring agent in the anchor rod drill hole is shown, and the unit is mm; dDrill 1The diameter of the anchor rod drill hole is shown, and the unit is mm;DrodThe diameter of the anchor rod is expressed in mm;
II, determining the ultimate elongation of the roadway roof anchor rod, and then calculating the energy consumed before the anchor rod generates ultimate deformation and is broken according to the ultimate elongation of the roadway roof anchor rod: firstly, acquiring the non-anchoring section length of the anchor rod and the elongation of the anchor rod, and substituting the acquired data into the following formula:
he1＝Lis not 1×μ1
In the formula, he1The ultimate elongation of the anchor rod is expressed in m; l isIs not 1The length of the non-anchoring section of the anchor rod is expressed in m; mu.s1Indicating the elongation of the anchor rod;
calculating the energy E consumed before the anchor rod generates the limit deformation and is broken according to the obtained limit elongation of the anchor rodg；
III, determining the length of an anchoring section of the roadway roof anchor cable: firstly, acquiring the compression coefficient of the used anchoring agent, the length of the anchoring agent in an anchor cable drilling hole, the diameter of the anchoring agent in the anchor cable drilling hole, the diameter of an anchor cable drilling hole and the diameter of an anchor cable; the obtained data are then substituted into the following formula:
in the formula: l is2The length of the anchorage section of the anchor cable is expressed in mm; k is a radical ofSolid 2Represents the compressibility of the anchoring agent in units of 1; l isSolid 2The length of the anchoring agent in the anchor cable drill hole is shown, and the unit is mm; dSolid 2The diameter of the anchoring agent in the anchor cable drill hole is shown, and the unit is mm; dDrill 2The diameter of the anchor cable drill hole is shown, and the unit is mm; dCableThe diameter of the anchor cable is expressed in mm;
IV, determining the ultimate elongation of the roadway roof anchor rod, and then calculating the energy consumed before the anchor cable is subjected to ultimate deformation and breakage according to the ultimate elongation of the roadway roof anchor cable: firstly, acquiring the non-anchoring section length of the anchor cable and the elongation of the anchor cable, and substituting the acquired data into the following formula:
he2＝Lis different from 2×μ2
In the formula, he2The ultimate elongation of the anchor rod is expressed in m; l isIs different from 2The length of a non-anchoring section of the anchor cable is expressed in m; mu.s2Representing the elongation of the anchor cable;
calculating the energy E consumed before the anchor cable is subjected to ultimate deformation and breakage according to the ultimate elongation of the anchor cableds；
And V, determining the energy limit value absorbed by the roof support system before the roof support system is broken according to the arrangement distance condition of roof anchor rods and anchor cables, wherein the energy limit value is as follows:
Edz＝ρ1×Eg×η1+ρ2×Eds×η2
in the formula, EdzRepresenting the limit of energy absorbed by the roof support system before rupture, in kJ/m2；ρ1The density of the anchor rod arranged on the top plate is shown, and the unit is root/m2；η1The unbalance coefficient of the energy consumption of the broken anchor rod is represented, and the value is 0.8; rho2The density of the anchor cables arranged on the top plate is shown, and the unit is root/m2；η2The imbalance coefficient of energy consumption when the anchor cable is broken is represented, and the value is 0.9;
VI, in a roof supporting system with rock burst, kinetic energy transferred to the roof rock mass and potential energy of sinking of the roof rock mass are transferred to the supporting system to become elastic energy of the supporting system, and in a critical state, the elastic energy borne by the supporting system is equal to the energy limit value of the supporting system, and the minimum speed v causing the roof supporting system to lose efficacy is calculated: the elastic energy borne by the supporting system is equal to the kinetic energy transferred into the roof rock mass and the sinking potential energy of the roof rock mass, and the specific formula is as follows:
in the formula, EdThe kinetic energy of the surrounding rock of the roadway in unit area after the rock burst occurs is expressed in kJ/m2；EhThe unit area potential energy applied to the supporting system by the roof coal blocks due to the downward sliding of the impact is expressed in unitkJ/m2(ii) a m represents the mass of the rock mass of the surrounding rock of the roadway participating in the impact process, and the unit is kg; v represents the impact velocity of the surrounding rock of the roadway, and the unit is m/s; s represents the cross-sectional area of the roadway participating in the impact process, and the unit is m2(ii) a g represents the gravity acceleration of the position where the roadway is located and the unit is m/s2(ii) a Δ h represents the distance of the top plate coal block sliding down due to impact, and the unit is m; wherein S, m, g and delta h are known data;
due to the critical state, the elastic energy to which the timbering system is subjected is equal to the energy limit of the timbering system, i.e. EdzE; the minimum velocity v that causes the roof support system to fail is given by:
thereby deriving a minimum velocity v that causes failure of the roof support system;
VII, evaluating the impact resistance of the roadway roof anchoring support: acquiring data of the mine earthquake energy level obtained by monitoring the mine, the historical mine earthquake energy level or the surrounding mine earthquake energy level, selecting the maximum rock mass vibration speed corresponding to the maximum mine earthquake energy level to compare with the minimum speed v obtained in the step VI, and if the maximum rock mass vibration speed is less than the minimum speed v, indicating that the impact resistance of the roadway roof anchoring support meets the required requirement; and if the maximum vibration speed of the rock mass is greater than or equal to the minimum speed v, the requirement on the impact resistance of the roadway roof anchoring support is not met, and at the moment, subsequent support reinforcement work is carried out according to the difference value between the maximum vibration speed of the rock mass and the minimum speed v.
2. The method for calculating the impact resistance of the roadway roof bolting support according to claim 1, wherein the energy E consumed before the anchor rod is subjected to ultimate deformation and breakage is calculated in the step IIgThe following formula:
Eg＝Ef1×he1×f1
in the formula: ef1Indicating the anchor rod producing ultimate deformation and breakageBreaking load in kN/root; h ise1The ultimate elongation of the anchor rod is expressed in m; f1 represents the elongation safety factor when the anchor rod is in tensile failure; wherein Ef1And f1 are known values.
3. The method for calculating the impact resistance of the roadway roof bolting support according to claim 1, wherein the energy E consumed before the anchor cable is subjected to ultimate deformation and breakage is calculated in the step IVdsThe following formula:
Eds＝Ef2×he2×f2
in the formula: ef2Representing the breaking load when the anchor cable is subjected to ultimate deformation and breaking, wherein the unit is kN/root; h ise2The ultimate elongation of the anchor cable is expressed in m; f2 represents the elongation safety factor when the anchor cable is damaged by stretching; wherein EdsAnd f2 are known values.
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| CN112903480A (en) * | 2021-01-15 | 2021-06-04 | 安徽理工大学 | Deep roadway anchor rod or anchor cable impact tensile damage judgment and control method thereof |
| CN113283054A (en) * | 2021-04-20 | 2021-08-20 | 中国科学院武汉岩土力学研究所 | Calculation method of anchor rod reinforcement effect and related equipment |
| CN115326601A (en) * | 2022-10-14 | 2022-11-11 | 中国矿业大学(北京) | Dynamic impact test and evaluation method for anchor net coupled supporting rock mass |
| CN115788573A (en) * | 2022-12-07 | 2023-03-14 | 华北科技学院 | Method for evaluating anti-vibration capability of roadway roof support system |
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| CN115788573A (en) * | 2022-12-07 | 2023-03-14 | 华北科技学院 | Method for evaluating anti-vibration capability of roadway roof support system |
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